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L.R. Bower In regard to the paper by J.B. Gwiazda, it makes a highly technical approach to show that the µ factor used by designers of lemniscate-guided roof supports has never really been confirmed as a maximum and assumes that convergence is vertical. Also, the paper does not appear to take into account deflection of structures, which occurs when the lemniscate and base members are fully loaded to their maximum stress level, nor the front to back line of the support in relation to differential roof to floor movements caused by strata movements under pressure. It is not unusual for differential movements to be slightly diagonal to the line of the support, particularly in faulted areas and on gradient faces. The paper also does not take into account consolidation of fines immediately above and below the support. Generally speaking, any differential movement is from face to waste and under these conditions the µ of 0.3, which appears to be an international standard, has worked in practice. However, if the face end of the support is lower than the waste end, then the µ of 0.3 can be considerably increased, giving rise to the damage mentioned in the paper. The ideal design should aim for a slightly forward bias in the lemniscate guide so that the last increment of setting is toward the face, tending to close any fissures that may have developed during the support advance cycle. The support should also be fitted with positive set valves to ensure that a high setting load density is attained to minimize bed separation. As far as powered supports are concerned, convergence is irresistible and all powered supports converge at their rated yield load. A similar principle can be applied to the differential roof to floor movements to drastically reduce the very high forces that would otherwise be applied to the lemniscate structures and pins and that, in turn, are transferred to the base arrangement and floor loading. Any differential movements are usually catered for by the 0.3 µ factor or deflection of structures in the lemniscate guide arrangement and consolidation of the floor. The floor loading, due to differential movement, is in addition to the support convergence load and requires additional bearing area to avoid possible floor penetration. Some seven years ago, Fletcher Sutcliffe Wild Ltd. (FSW) introduced a lemniscate-guided shield support where the lemniscate linkage is connected to the roof bar through two horizontally converging rams to allow differential movement to take place above a given rated figure. This is a known force and can be guarded against, whereas with rigid connections the forces, as yet, are unconfirmed. By careful design, a horizontal force in excess of 6 MN (60 tons) opposes differential movements for a total ram loading of only 2.5 MN (25 tons), or 1.25 MN (12% tons) each. This principle can considerably reduce the length and weight of the support in comparison with a rigid pin-type structure ; also, the yield load rating can be increased without affecting the lemniscate forces. The graph shows the tensile and compressive forces in a lemniscate linkage of a support with and without hydrostore. These forces react into both the roof beam and base members and, as can be seen from the support height to linkage load graph, a considerable reduction in these reactions is gained by the use of the FSW patented hydrostore system. Floor loading is considerably reduced under maximum µ conditions, and by allowing the roof bar to move with the strata, some degree of improvement to strata control is achieved in line with the assumptions in the paper. In practice, these movements have only been in the region of a few millimeters, which, in turn, reflects on the improvements to strata control by the addition of positive set valves. Supports to this design of both 450- and 280-t (496-and 309-st) rating have been successfully used in the United Kingdom for several years, negotiating many faulted areas without one single reported need for repair or maintenance. This includes supports left unattended during the year-long strike, proving the reliability of the system.

Jan 8, 1986

By Mark A. Krall, Robert L. Geddes, James C. Frost

Introduction The Conda Partnership's Dry Valley phosphate mine is a thinly bedded, multiple seam open-pit mining operation where selective mining techniques are used to recover phosphatic shales. The mining methods used are truck/shovel and scraper/dozer operations. Ore is shipped 32 km (20 miles) by rail to a beneficiation facility. The ore is upgraded by washing and calcining. The mine and beneficiation complexes are owned by the Conda partnership. It is a joint venture between Beker Industries Corp., of Greenwich, CT, and Western Co-Operative Fertilizers (US) Inc., of Alberta, Canada. The Partnership operates as a separate entity of the two partners. The Dry Valley mine is located 48 km (30 miles) northeast of Soda Springs in Caribou County in southeastern Idaho. The mine is situated on the Caribou National Forest. Mining operations take place between 2 and 2.4 km (6400 and 7900 ft) in elevation. It is accessible partly by 32 km (20 miles) of paved roads and 16 km (10 miles) of dirt roads. The winters are long and severe, and the summers are short and mild. This article describes the history, geology, exploration, mining, and reclamation that makes this mine Idaho's largest producing mine and the western US' leading phosphate producer. History and production In the mid-1950s, Western Fertilizers of Salt Lake City, UT, drove an exploratory drift in Maybe Canyon. A large bulk sample of phosphatic shales was analyzed for phosphate content and processing characteristics. No large scale mining or processing operations were undertaken. In the late 1950s, the Dry Valley property was sold to Central Farmers of Chicago, IL. No major operations took place. In 1964, Central Farmers sold the property to El Paso Products Co. of Odessa, TX. El Paso Products supervised the mining operations of Wells Cargo Mining Co. from 1965 through 1967. During this time, El Paso Products built a beneficiation facility and a fertilizer complex in Conda. A 32-km (20-mile) railroad was also constructed from the mine to this facility. From 1968 through 1972, the mine was shut down due to a depressed fertilizer market. In 1972, El Paso products sold its ore reserves, beneficiation plant, and fertilizer complex to Beker Industries Corp. In 1979, Beker Industries sold 50% of its ore reserves and 50% of its beneficiation plant to Western Co-Operative Fertilizers (US) Inc., of Alberta Canada, forming the Conda Partnership. It has operated the mine and beneficiation plant since January 1979. From the mid-1950s to the mid-1960s, no substantial production took place. From 1965 to 1967, El Paso Products stripped 3 Mm3 (4 million cu yd) and mined 2.3 Mt (2.5 million st). From 1972 through 1983, 50 Mm3 (66 million cu yd) were stripped and 18 Mt (20 mil¬lion st) were mined. Geology The Wells Formation forms high ridges and hillsides in the Dry Valley area. It is best exposed along the west face of Dry Ridge. It forms the imposing wall on the east side of Dry Valley. The formation is divided into two members. The lower member, about 213 m (700 ft) thick, is dominantly thin to medium-bedded limestone and silty limestone. It contains nodules and stringers of chert and minor sandstone. The upper member is composed principally of thick-bedded to massive cross-bedded, light-gray to orange-yellow, fine grained sandstone. There is some interbedded brown to light-gray limestone. This member varies from 369 to 457 m (1300 to 1500 ft). Recent investigations indicate that the upper Wells is of Permian age. Under some conditions, the Wells may be water-bearing. Otherwise, it has no apparent economic significance. Grandeur Member (Park City Formation) Overlying the Wells Formation is a distinctive light-gray to white dolomitic fossiliferous limestone. This unit has been identified by the US Geological Survey (USGS) as the Grandeur Tongue Member of the Park City Formation. This member is sometimes absent due to its contact with the Meade Peak Member of the Phosphoria Formation. It is easily detectable by its color, hardness, and fetid odor. Phosphoria Formation The Phosphoria Formation of Permian age was named from Phosphoria Gulch, Bear Lake County. The formation has been studied extensively and developed for its economically valuable phosphate reserves.

Jan 11, 1985

By Edward Burris, Terry Gorsuch, Joseph M. Hans

INTRODUCTION In any large earth-moving operation, good planning and management are necessary to complete the operational tasks promptly and successfully. When an earth moving operation is complicated by radioactive contaminants, normal earth moving techniques and procedures must be modified. Any planning and management, therefore, must include the radiological aspects of the operation. It was found that the radiological aspects dominated most of the planning and management activities and were extended to all facets of the decontamination work at the former Shiprock uranium millsite. These planning aspects are discussed and their use to develop a work plan is described. The management aspects are discussed and their use to establish a management structure are also presented. PLANNING Some method of procedure, formulated beforehand, was necessary to govern the decontamination work at the former Shiprock uranium millsite. This procedure was expressed in the form of a work plan which served several listed purposes. 1. It defined the work to be done and the sequence it would follow. 2. It was used as a yardstick to measure progress. 3. It was used to assign organizational responsibilities. Several factors were considered to aid in the development of the plan. These factors are discussed below: Goals It was established that radiation exposure was occurring to persons working at the millsite, and in an around the community of Shiprock, from airborne radioactive mill wastes and radon-222 exhaling from the tailings piles. The goal set for the decontamination work was to reduce on-site exposures to levels acceptable for the millsite occupants. The attainment of this goal would also have a substantial impact in reducing off-site exposures. The objectives necessary to achieve the goal were consolidation and containment of the wastes. The former objective implies decontamination of the millsite and environs, and the later implies stabilization of the wastes. In practice, a total and complete decontamination of the millsite and contaminated environs would be very difficult and costly. The costs for decontaminating them could be high enough that an alternative method might be more cost-effective for reducing human exposures (i.e, move the affected people away from the source). The interim guide "Radiological Criteria for the Decontamination of Inactive Uranium Millsites" was used for the decontamination criteria (EPA 74). Briefly, the criteria state that off-pile decontamination should be effective enough to reduce the net above ground exposure rate to less than 10 [u]R/hr for unrestricted use of the affected area. When decontamination cannot readily be achieved, the exposure rate levels could be relaxed to 40 [u]R/hr; however, the affected area has to be restricted. The second objective, waste containment, means isolating the wastes from the biosphere. Since no method of containment was available at the beginning of the millsite decontamination effort, temporary containment (interim stabilzation) became the objective. The tailings pile and decontamination wastes would be covered with clean fill. The interim stabilization should last from 5 to 10 years until the final disposition of the wastes will occur. The goal, therefore, would be achieved by decontaminating the off-pile areas to less than 10 uR/hr where practical. The decontamination wastes would be used to plate the surface of the tailings pile and would be covered with clean fill. Radiological Survey The radiologial survey is the key factor for planning a decontamination activity. The survey should delineate the spread and depth of the contaminants relative to the decontamination criteria. Surface wastes, in general, can be evaluated for spread and depth with reasonable radiation survey equipment. Subsurface wastes on the other hand can be missed entirely, as happened during the radiation survey at the Shiprock site, although numerous exploration holes were bored and dug. The survey results can be used to define areas that may not be amenable to decontamination because of complications or safety reasons. For example, no decontamination of the bluff base was to be attempted because of the possibility the bluff might collapse on the personnel and equipment. Contaminated bottoms of decant ponds on the flood plain were not removed because they would be slurried by ground water. Slurry removal was deemed inefficient because the contaminants would be scattered and no equipment was available for its transport. In summary, the radiological survey defines the boundaries of the decontamination work and provides

Jan 1, 1981

By Gregory R. Wessel

Sulfur is a nonmetallic element widely distributed in nature and of great physical and economic importance. It is the 14th most abundant element in the Earth's crust (0.06%) and an important constituent of animal and plant life. Sulfur has been known and used since ancient times for a number of medicinal and industrial uses. At present, most sulfur is used to generate sulfuric acid that is used in a wide variety of industrial processes, particularly the production of fertilizer. Because of this, sulfuric acid (and hence sulfur) consumption is often regarded as a good index of a nation's industrial development. In the past, sulfur was mined from surface occurrences in several geologic environments, and was used in relatively small amounts. With time, the uses of sulfur and sulfuric acid expanded, as has the need for larger quantities of these commodities. Sulfur is now mined from both surface and underground deposits, and is recovered as a byproduct from a number of industrial processes. Despite valiant efforts and years of work by sulfur explorationists and others, many aspects of sulfur mineralization remain controversial. Almost as controversial is the spelling of the word sulfur. The English spelling sulphur commonly is used outside America and in the American sulfur mining industry, but sulfur is the correct American spelling as approved by the American Chemical Society, the American Geological Institute, and many others. Those new to the American sulfur industry often find it puzzling to be reprimanded for using the correct American spelling. Sulfur resources are abundant and widespread, but the extent to which they can be classified as reserves is constrained by pre- vailing prices and extraction technologies. At present, sulfur can be economically mined from very few deposits. The sulfur industry is roughly divisible into two sectors: voluntary (or discretionary) and involuntary (or nondiscretionary). In voluntary production, the mining of sulfur or pyrites is the sole objective, and the recovery of the resource is as complete as economic conditions will allow. During involuntary production, sulfur or sulfuric acid (termed recovered sulfur) are produced as byproducts, and the quantity of the output is dictated by the demand for the primary product. Voluntary sulfur now accounts for only about 35% of the elemental sulfur produced worldwide, and most inves- tigators believe that voluntary sulfur will be less important in the future. Sulfur sources and products are described as follows (after Barker, 1983): Sulfur Sources: Combined sulfur-sulfur that occurs in nature combined with other elements, commonly referring to sulfides and sulfates. Cupriferous pyrites-pyrite containing minor amounts of cop- per sulfides. Hydrogen sulfide-a toxic gas that occurs in petroleum and natural gas. Involuntary sulfur-sulfur produced as a byproduct in response to legislative or process mandates. Native sulfur-naturally occurring elemental sulfur. Nonferrous metal sulfides-opper, lead, zinc, nickel, and molybdenum sulfides that are processed for their metal content. Organic sulfur complex organic sulfur compounds that occur in petroleum, coal, oil shale, and tar sands. Pyrites-iron sulfide minerals that include pyrite, marcasite, and pyrrhotite. Sulfate sulfur-sulfur contained in anhydrite and gypsum. Voluntary sulfur-sulfur produced in response to market demand. Basic Sulfur Products: Acid sludge-contaminated sulfuric acid usually returned to acid plants for reconstitution. Brimstone-synonymous with crude sulfur. Bright sulfur-crude sulfur free of discoloring impurities and bright yellow in color. Broken sulfur-solid crude sulfur crushed to -8 in. Byproduct sulfuric acid-sulfuric acid produced as a byproduct of a metallurgical or industrial process, generally relating to combined sulfur sources. Crude sulfur-commercial nomenclature for elemental sulfur. Dark sulfur-crude sulfur discolored by minor quantities of hydrocarbons, ranging up to 0.3% carbon content. Elemental sulfur-processed sulfur in the elemental form produced from native sulfur or combined sulfur sources, generally with a minimum sulfur content of 99.5%. Formed sulfur-elemental sulfur cast or pressed into particular shapes to enhance handling and to suppress dust generation and moisture retention. Frasch sulfur-elemental sulfur produced from native sulfur sources by the Frasch mining process. Liquid sulfur-synonymous with molten sulfur. Liquid sulfur dioxide-purified sulfur dioxide compressed to the liquid phase. Molten sulfur-crude sulfur in the molten state. Prilled sulfur-solid crude sulfur in the form of pellets produced by cooling molten sulfur in air or water. Recovered sulfur-elemental sulfur produced from combined sulfur sources (including byproduct hydrogen sulfide, but sometimes referring only to sulfur from fossil fuels) by any method. Slated sulfur-solid crude sulfur in the form of slate-like lumps, produced by allowing molten sulfur to solidify on a moving belt. Specialty sulfur-prepared or refined grades of elemental sulfur that include amorphous, colloidal, flowers, precipitated, wettable, flour, and paste sulfur. Sulfur ore-unprocessed ore containing native sulfur.

Jan 1, 1994

By L. H. Munson, D. E. Hadlock, L. F. Munson, R. L. Gilchrist, P. D. Robinson

All uranium mills are required to perform sampling and analysis for radioactive particulates in their gaseous effluent streams and in the environment. Pacific Northwest Laboratory was requested by the U.S. Nuclear Regulatory Commission (NRC) to provide technical assistance to them for their Uranium Mill Health Physics Appraisal Program. In conducting appraisals, air sampling methods used at NRC-licensed mills were reviewed and several deficiencies noted. This paper includes only environmental and effluent particulate sampling although much of the information is applicable to both in-plant and environmental samples. First, the components of a proper sampling program are discussed: program objectives, program design, sampler design, analyses, quality assurance, and data handling. Then the specific deficiencies, or the "pitfalls" from the first 8 mill appraisals are discussed. The first consideration in establishing an air sampling program is defining the objectives of the program. What is air sampling suppose to accomplish? Many of the deficiencies we have observed have resulted because the desired objectives were not clearly established in the minds of the radiation safety staff. PROGRAM OBJECTIVES An environmental air sampling program ought to fulfill the following seven objectives. The first is to: 1) [demonstrate regulatory compliance]. Although a goal of most programs, regulatory compliance, is not well understood. One has not only to comply with the conditions of the source materials licensee, but one must also demonstrate compliance with 10CFR20 and 40CFR190. For example, 10CFR20.106 states: "A licensee shall not possess, use, or transfer licensed material so as to release to an unrestricted area radioactive material in concentrations which exceed the limits specified in Appendix B, Table II of this part .... For purposes of this section, concentrations may be averaged over a period not greater than one year." Even if a mill's license does not require sampling at the site boundary of maximum concentration, a sample may be necessary to demonstrate compliance with 10CFR20. Most mill personnel are painfully familiar with 40CFRl90.10, which states: "Operations.... shall be conducted in such a manner as to provide reasonable assurance that: (a) The annual dose equivalent does not exceed 25 millirems to the whole body.... of any member of the public as the result of exposures to planned discharges of radioactive materials, radon and its daughters excepted... from uranium fuel cycle operations..." This means a licensee's sampling program must give "reasonable assurance" that the member of the general public receiving in the most exposure gets no more than 25 millirems per year. The sampling program necessary to provide that assurance may or may not be a license requirement. However, merely meeting the license requirements and the explicit regulatory requirements does not necessariarly ensure an adequate effluent and environmental air sampling program. The second objective of the environmental air sampling program, is to 2) [identify the source(s) of contaminants]. This will include not only the routine program, but special sampling for verification of sources and nonsources. Only after sampling can a mill operator be assured that roof vents, laboratory hoods, and other localized ventilation systems are not making a significant contribution to environmental releases. An environmental sampling program should also allow the mill operator to fulfill the third objective, to 3) [estimate exposures]. Even before 40CFR190, a sampling program should have provided the mill operator with the information necessary to determine the dose to the "fence post" person, or at least to determine if doses were well below the 10CFR20 limits previously allowed. The program should 4) [detect and measure unplanned releases]. If there is a fire, a scrubber failure, or if a drum of yellowcake breaks open, measured releases will almost always be lower than conservative estimates. Whether or not a system to provide sampling during accidents is needed is almost always a cost-benefit decision. In general, uranium operations do not sample just in case an accident may occur. Yet they may decide on continuous air sampling in lieu of intermittant sampling partially because of the potential for accidents. Another objective of air sampling is 5) [to provide information on the effectiveness of control systems]. This is always a concern with new or modified equipment and may dictate sampling frequency in other situations as well. For instance, if a small leak in a bag filter cannot be detected by other means, then more frequent stack sampling may be indicated. A routine effluent and environmental monitoring program should also fulfill the sixth objective,

Jan 1, 1981

By Aladar B. Dory

INTRODUCTION Practical experience in mines with known presence of radon daughters in the mine atmosphere in Canada and elsewhere shows that a very high concentration builds up in an unventilated dead end heading. As Holaday et al1 observed, even a minimal air movement results in a drastic reduction in radon daughter concentration. It is therefore obvious that the main objective of radon daughter control in the working environment is to design the ventilation system providing an optimized flow of fresh air into the workplace, resulting in acceptable climatic conditions and achieving radon daughter concentrations resulting in exposures as low as reasonably achievable. BASIC OBJECTIVES Large mining companies, having extensive material resources and professional expertise, have utilized elaborate electrical modelling in the design of mine ventilation systems as early as 1950 (coal mining industry in Europe) and with the advance of computer modelling techniques, their utilization in ventilation systems design is on the increase. Unfortunately, these methods are usually not available to small mining companies and even the large companies might not achieve the fullest benefit from utilizing them, if proper limiting factors are not considered in the modelling. When an evaluation of a ventilation system of a mine is undertaken in literature, a measure of the amount of air supplied underground per one ton of ore mined is used as an indicator of the efficiency of the ventilation system. Yet, even the greatest amount of air forced into the mine might not result in an acceptable working environment if a proper distribution of this air into individual working places is not achieved. The volume and the age of the air are probably the two most important factors in achieving acceptable radon daughter concentrations in the workplace, but other factors also have to be considered. DIRECTOR MINE - ALCAN, NEWFOUNDLAND FLUORSPAR WORKS ST. LAWRENCE, NEWFOUNDLAND, CANADA Ventilation To illustrate the effects of the design of the ventilation system on the control of radon daughter concentration, let us review the gradual development of the ventilation system of this mine from the earlier years of its development up until its final years of operation. This mine, located near the community of St. Lawrence on the south coast of Burin Peninsula was developed in the late thirties and reached full production by 1942. Unfortunately as was customary at that time, the only source of ventilation was a natural draft. The mine was extremely wet, and no significant attention was initially given to possible health effects of dust. It was not until the mid-fifties, when a number of cases of silicosis had surfaced, that de Villiers and Windish2 observed a significant increase of lung cancer incidence among the miners in comparison to its incidence among the general population of Newfoundland. Suspicions regarding radiation as a cause of the lung cancer were expressed, but it was only in surveys taken in late 1959 and early 1960 that Windish3 and Little4 established the presence of radon daughters in the mine atmosphere in very high concentrations. Windish, de Villiers and Hurley suggested that the most likely source of the radon in the mine was the mine water which dissolved radon during its passage through the granitic country rock in the surrounding geological area. This conclusion was confirmed by analyses of water from various areas of the mine by the Atomic Energy Canada Limited laboratories. The radon values in the samples varied from 4,240 to 12,850 pCi/L5. Following the discovery of the presence of radon daughters in the mine, the company took speedy action to install mechanical ventilation for the mine. The system was not designed as a total unit, but fans were installed rather on a trial and error basis. The basic system installation began in March 1960 and was completed by 1962. It remained basically unchanged with only minor modifications until August 1973 when a wholly new, redesigned ventilation system was implemented. A schematic section of the mine and its ventilation system for the period prior to March 1960 is given in Figure "A", for the period 1960-1973 in Figure "B", and for the period after August 1973 in Figure "C". The ventilation system prior to 1960 is not known. All workings of the mine were ventilated only by natural ventilation. If any measurements of airflows at different or any times of the year ever existed, no records have been preserved. The very minimal natural ventilation was augmented by "blowing" air from compressed air supply lines and exhaust air from drills. It is known that the compressor capacities of the mine were limited and therefore no significant air movement was probably created by the "blowing".

Jan 1, 1981

By W. A. Haley

The use of diesel engines in underground hard-rock mines dates back to the late 1940s. For the first several years, they were used only occasionally, being limited to a few metal mines that experimented with crawler¬mounted front-end loaders, tractor-trailer hauling units, a few tractors for drill-compressor mounts, and utility cleanup machines. By the mid-1950s, track loaders had become commonplace in limestone mines and uranium mines on the Colorado Plateau in the United States, as well as in Canada. Use of crawler-mounted tractors as drill and compressor mounts also increased. By the end of the 1950s, rubber-tired loaders and some haulers began to replace the track-type machines and rail-mounted cars that had been in use. About 1960, the rubber-tired machines brought about a new era of underground mining mobility and flexibility, centered on a method commonly known as "trackless mining." Ultimately, many of the underground rail-type systems for loading and hauling were replaced by the trackless mining technique. ECONOMIC CONSIDERATIONS The size and nature of mineral deposits, plus ground control techniques, historically had dictated small open¬ings to the surface from many underground mines. The small mine openings led to the development of special rubber-tired loaders and haulers designed specifically for access through the small openings. However, some mines, particularly those in massive mineral deposits, are able to excavate and maintain very large openings, and some use modified room-and-pillar systems. With the large mine openings, the use of larger, more produc¬tive equipment such as that commonly found in surface mining becomes economical. In fact, productivity gen¬erally increases at a more pronounced rate than machine size increases because many of the larger machines were designed for heavy-duty shot-rock applications in surface mines and construction sites where the handling of blasted rock is common. Table 1 can be used as a very Table 1. General Productivity Comparison for Conventional Machines In Underground Use (Shot-Rock Conditions) 2.3 m3 (3 cu yd) 4.6 m3 (6 cu yd) Loader Loader Expected Surface 230 t/h 540 t/h Production (250 stph) (600 stph) Expected Underground 90 t/h 270 t/h Production (100 stph) (300 stph) Expected Total Efficiency: Surface 40%-60% 50%75% Underground 25%.-40% 30%50% Expected Useful Machine 8000 hr 12,000 hr Life Before Replacement general comparison of the production and efficiency between small and large machines. Combining greater productivity often inherent in larger machines, with reduced downtime resulting from using fully developed machines with fast parts and service backup, some mine operators have been able to reduce material handling costs appreciably while reduc¬ing manpower requirements for operators and main¬tenance men. Large mine openings increase the amount of rock that must be handled in the development work, and they sometimes increase the dilution in stopes or rooms, de¬pending upon the dimensions of the ore zone. Providing adequate space for the unrestricted operation of large surface mining machines could, therefore, lead to more waste segregation and handling costs. It could also cause greater ore dilution that would result in a lower grade of ore being delivered to the processing plant. The tradeoffs between opposing cost factors must be reconciled and balanced to achieve the best overall cost of the crude ore, concentrates, or product. EQUIPMENT MODIFICATIONS Loaders and haulers designed for surface mining are seldom used underground in their standard con¬figurations without some modifications. If done, the modifications generally are made by the equipment dealer and/or the user, and the modifications usually include one or more of the following items: 1) The exhaust stack is lowered, and its direction is changed. Usually, it is repositioned horizontally to the rear, or it is fed into the engine fan to diffuse the exhaust gases. 2) The operator's position is lowered by either lowering the seat or changing the seat to a side mount. 3) The operator controls are adjusted to fit the new operator position. 4) Other components, such as the radiator and loader tower, are lowered. 5) Special bumper guards are mounted at the base of the radiator area. 6) An exhaust conditioner is mounted and con¬nected, using either a catalytic or a water-type condi¬tioner, or both. This usually is controlled by the safety and health regulatory authority having jurisdiction. 7) The positions of other components are rearranged

Jan 1, 1982

By Ivor Hawkes

INTRODUCTION High energy breaking is an alternative to using ex¬plosives in underground secondary breaking operations. It also is a means of upgrading conventional hand-held breakers, manual sledge-hammer breaking, and scaling bar operations. Major areas of application are in sec¬ondary breaking over grizzlies and at drawpoints. Other applications include breaking down ripping lips in longwall seam mining, scaling in stopes and rooms, general demolition work, and roadway maintenance. There is considerable interest in high-energy impact breakers for use in primary ore breaking, but, as of 1977, all such applications have been only experimental (duToit, 1973; Joughin, 1976; Wayment and Grantmyre, 1976). EQUIPMENT Essentially, a high-energy impact hammer is a boom¬mounted pneumatically or hydraulically actuated breaker. The machine basically consists of a piston that oscillates in a housing and impacts the end of a tool or moil thrust against the rock. The force applied to the rock primarily depends upon the impact energy of the piston-the higher the impact or blow energy, the greater the force and, thus, the greater the rock break¬age. Among drill and breaker designers, a common expression for blow energy is "force of blow." Hand-held breakers are limited to blow energies of about 140 J (100 ft-lb), because the operator is unable to handle heavier machines efficiently or to absorb the recoil energy resulting from higher blow energies. How¬ever, these restrictions do not apply to boom mounted breakers; machines with blow energies on the order of 4000 J (3000 ft-lb) and higher are available commer¬cially for underground use. There is considerable evi¬dence to show that increasing the blow energy also in¬creases the efficiency of the breaking operation; i.e., more rock is broken per unit of energy expended (Grantmyre and Hawkes, 1975). Thus, there is a trend to higher blow-energy machines, particularly where high¬strength rocks are to be broken. In relation to rock breaking, the blow rate of boom¬mounted impact breakers is not as important as it is for rock drills. This is because the breaker must be moved over the work surface between blows. The blow rate is governed eventually by the power supply, and typical blow rates range between 200 and 600 blows per minute. As a general rule, light blow-energy machines have higher blow rates than heavier machines. Table 1 lists most of the boom-mounted impact breakers that were available commercially during 1977, and it gives details of the blow energies and machine weights. Restrictions are placed on the blow energy by the machine weight and size, and by the strength of the boom. Typically, boom-mounted impact hammers have a blow-energy to mass ratio of about 1.5, with lower values for lighter machines and higher values for heavier machines. In addition to supporting the hammer weight, the boom also has to absorb the recoil energy of the blow, which can be on the order of 1400 J (1000 ft-lb) for large hammers operating in a horizontal mode. Interesting exceptions to the general run of impactors are the Joy HEFTI hydraulic hammers. In these machines, the piston impacts onto a fluid cushion that is positioned between the piston and the impact tool. This approach allows very high piston velocities, over 30 m/s (100 fps), to be used without the risk of break¬ing the piston or impact tool. Steel on steel impacts must be limited to impact velocities of about 10 m/s (35 fps) due to the high impact stresses generated; thus, increased blow energies can be achieved only by increas¬ing the piston size. The Joy 514 HEFTI®, listed in Table 1, has a blow energy of 27 100 J (20,000 ft-lb), but, as of 1977, the machine has been used underground only on an experimental basis. Using a fluid cushion between the piston and the impact tool allows the use of light pistons, reducing the overall machine weight. The recoil energy, which must be absorbed by the boom for a given blow energy, is directly proportional to the piston to machine mass ratio, and operating with light pistons provides an addi¬tional benefit in reducing the requisite boom size. Both pneumatic and hydraulic hammers are avail¬able commercially. Although hydraulic hammers are a relatively recent development, they already outnumber the pneumatic machines in use. There are many reasons

Jan 1, 1982

By M. Kalin, C. Steinberg

We have worked on several flooded pits from coal-mining activities in the former East Germany, as well as ones associated with hard- rock mining, including the B-zone pit discussed in the above technical paper. We found the paper to be a useful summary, but, unfortunately, it failed to give an adequate comparison of the physical limnology of the flooded pits, which is an essential component. While the title suggests that the primary focus of the review is physical limnology, it appears that it is essentially pit-lake chemistry being presented. Physical limnology requires that factors such as fetch, latitude, light penetration, relation to ground water table, methods of flooding and the physical shape of the pits be defined. These physical aspects of a pit interact with the chemical and biological processes taking place in it, all of which contribute to the character of a water body. Few of these physical aspects are presented, however. The conclusion that the authors reach suggests that meromixis may be a condition that would serve as an effective containment mechanism for contaminants in a pit. Although this may be desirable, such limnological conditions are not clearly supported by the data presented for any of the pits. These data should be summarized to facilitate comparison between the same structural units of the pit water - the epi- and metalimnion for example. The thermocline depth is a reflection of the physical forces mixing the water body, and pit dimensions affect these forces. Due to the use of different scales in Figs. 2 through 5, it is difficult to determine whether the thermocline is at the expected depth, because the fetch is not given. Moreover, the status of a water body cannot be determined unless measurements cover a period of at least one year, and depth profiles are completed to represent the entire depth of the pit. This shortcoming is most notable in the case of the Berkeley pit, where data are given for depths of only 20 and 35 m (66 and 115 ft), although the pit is reported to be 242 m (794 ft) deep. Limnological data to define the status of the pit water have to be collected at regular intervals, for the same parameters. The authors present temperature measurements for 1-m (3.3-ft) intervals, but fail to use that interval for other parameters, such as dissolved oxygen or, in some cases, for contaminant concentrations. Furthermore, the profiles for the deepest part of the pit display only part of the picture, because pits are rarely conical. Profiles can be considered to represent the status of a water body only after other stations in the pit have been monitored regularly and the consistency is determined. For example, fresh water, which can enter a pit at any depth, would interfere with the proposed meromictic conditions. Similarly, organic material at the bottom of a pit, such as the fish-waste deposited in the Gunnar pit, contribute to oxygen consumption. Oxygen depletion alone is not indicative of meromixis. It is interesting to note that the Dpit arsenic concentrations could possibly be slightly higher than the B-zone pit concentrations at depth, although this is difficult to determine accurately when a log scale is used for the D-pit and not for the B-zone pit. In our investigations, we noted arsenic removal in the B-zone pit bottom water, which was due to the formation of particles that are relegated to the newly forming sediment in the bottom of the pit. Particle-carrying contaminants form due to a combination of geochemical and biological factors and TSS contributed from erosion of the upper parts of the pit walls, whereas the settling out of particles from the water column is controlled by the physical conditions or turn over, for example. during ice cover in the B-zone pit. Although meromictic conditions for flooded pits may be desirable at decommissioning, this would depend largely on the physical conditions of the pit, because, under no circumstances, would this water be of desirable ground-water quality. Under meromictic conditions, acidity, if an environmental issue, may be reduced by microbial acid-neutralizing activity, and several heavy metals may form more or less stable sulphitic compounds. These may stay suspended in the water if conditions are such that they are not relegated to the sediments, i.e., in the absence of turnover. These processes do not take place in meromictic conditions only, but meromixis does require autochthonous and/or allochthonous organic substrate supplies, which are generated under aerobic conditions. Specific limnological (biological, chemical and physical) features of the pit lake under consideration have to be defined, such that water quality parameters can be predicted, and the objectives of the decommissioning activities, environ-

Jan 1, 1999

By Wade E. Cooper

INTRODUCTION How accurate are past and present employee radondaughter exposure records of underground uranium miners employed in the United States? This often-debated question is essential for future substantiation of safe exposure limits. An apparent discrepancy between company-reported exposures and Mining Enforcement and Safety Administration (MESA) projected exposures was detected in 1977. For these reasons a need for an updated comparison of these exposure data was indicated. This paper gives some of the conclusions of the earlier study and compares more recent exposure records compiled by the Atomic Industrial Forum, Inc., with projected exposures based on sampling by Federal mine inspectors. EARLIER STUDY In its 1977 Annual Report (U.S. Department of the Interior, 1978), MSHA's predecessor, the Mining Enforcement and Safety Administration (MESA), reported that there was "an apparent discrepancy between Federal inspection results and company records." Both company records and MESA's projections from samples taken during routine Federal inspections indicated reductions in the average exposure of underground uranium miners from 1975 to 1977, but the MESA projections were over 4 times higher than the company-reported averages. This apparent discrepancy however, was based on a comparison of exposure data reported for all U.S. underground uranium miners. This projection more closely represented the average exposure of U.S. underground uranium mine production workers who worked 1,500 hours or more during the year. Exposures of such workers are reported each year by the Atomic Industrial Forum, Inc. (AIF) in summaries of exposure data reported to the AIF by uranium mining companies throughout the United States. (The AIF exposure summary for 1979 appears as tables A-1 and A-2 in the appendix of this paper.) Assuming that the average exposure for each exposure range category is the midpoint of each exposure range category, table 1 compares the estimated average exposures for U.S. underground uranium mine production workers who worked underground 1,500 hours or more each year in 1975 through 1977 with the exposures projected by MESA for those years. [Table 1. - Average Exposure and Projected Average Exposure for U.S. Underground Mine Production Workers Who Worked Underground 1,500 Hours or More During the Year. Company, MESA?' Reported- Projected Year (WLM) (WLM) 1975 1.59 5.68 1976 1.84 4.64 1977 1.68 4.08 1 Atomic Industrial Forum, 1976, 1977, 1978. 2 U.S. Dept. of the Interior, 1978.] Table 1 indicates that, even after adjustment to ensure better comparability an apparent discrepancy between Federal inspection results and company reported exposures for 1975-1977 exists; however, the apparent discrepancy diminished over the 3 years. Slade, 1977, explained some of the discrepancy between company records and MESA projections of miners' average radon-daughter exposures as follows: 1) Concentrations of radon daughters in some work areas can vary greatly during any one day. A variation from 0.3 WL to 17.0 WL has been measured in the same stope on the same day. 2) Seemingly simple abatement problems indicated by the regular Federal and State inspections were solved simply by manipulating the mine ventilation. 3) The methods used by mine operators to compute cumulative exposures were such that high radiation readings were seldom or never reflected in the records. For example, a work area sampled on Monday indicated a radon-daughter level equal to 0.2 WL and this was recorded. It was sampled again on Wednesday of the following week and the level was 2.2 WL. The miners were withdrawn or told to fix the ventilation, and when this was accomplished the area was sampled and found to be at 0.2 WL again. Although the miners could have been working in the higher concentration up to 6 days, this reading might never be reflected in their records. If it was recorded, only a fraction of the day on which it was discovered would be entered into the cumulative exposure calculation (time-weighted average). 4) Some of the mines visited used a mine average radiation concentration, and every employee working underground was given the same exposure per unit of time spent underground. As a result of the 1977 study, more stringent sampling and recordkeeping standards were proposed and public hearings held in 1977. The resulting new and revised health standards on radon-daughter sampling and exposure recordkeeping became effective August 30, 1979 (Mine Safety and Health Administration, 1979). Prior to these new regulations, radondaughter sampling requirements were on an "as often as necessary" basis (Code of Federal Regulations, 1978). The new regulations required practically all active work areas in underground mines to be sampled at least once every 2 weeks, with many areas requiring weekly sampling. They also required calendar-year exposure records of all underground

Jan 1, 1981

By A. Dale Wunderlich

With the price of precious metals at an 18-year low, every ounce of metal produced is important. The theft of metals from mining and refining sites can mean the diffrence between profit and loss for many mining companies. Low metal prices do not reduce the potential for the theft of precious metals. History has shown that the price of gold has little to do with the desire for employees to steal precious or base metals. There is actually evidence that the theft of precious metals increases when the price of this commodity goes down. Several of the major precious metal thefts in the past year took place at silver mines when the price of silver was less than 16 cents/g ($5/oz). How does the lowest gold price in 18 years affect the need for security at precious metals properties? There is no short answer to this question. One reason is because the exposure to theft of precious metals is unique to each property. This makes it important that each property be evaluated individually. More than 95% of all precious metals thefts can be attributed to those working at the mine site. So preventing employee theft is the primary concern. One consideration is the location of the property. Gold selling at any price is still an attractive commodity in countries where the employees are making between US$400 and US$600 a month. It is not uncommon for employees at mines in countries where low wages are the norm to consider the value of a gram or two of gold to be a significant amount of money. A gram or two of gold a day may not seem like much. But if 15 employees steal two grams a day, that equates to a significant amount of money during a year. The type of property where the precious metals product is being recovered is also important. For example, a property with a gravity circuit is more likely to suffer from the theft of gold product than a property where all gold is finely disseminated and the only gold seen in the ore body is through a microscope. Gravity circuits increase an operation's exposure to theft because the grinding circuit that is associated with a gravity circuit often becomes a giant concentrator. Areas such as the bottom of grinding-mill pump boxes, cyclone-feed-pump clean out traps and the sumps often become locations where precious metals concentrate (Figs. 1 and 2). Muck concentrations in these locations can be as high as 25% to 40% of gold or silver. Not long ago, muck was removed from a barren-solution sump at a Merrill Crowe circuit that had concentrated to more than 40% gold. At a milling site in the Pacific Rim, residents of the community adjacent to the mine learned about the value of the concentrates in the sump under the ball mill and committed an armed rob¬bery. While several of their co-conspirators held the em¬ployees at bay with machetes, the others emptied the contents of the sump into buckets and removed it from the site. Armed robbery is not as common as employee theft. However, while this article was being written, an armed robbery occurred at a gold property in Central America. Armed perpetrators took as hostages the night shift employees at a process plant and used cutting torches that were on site to cut into the high-security and gold-storage areas. The perpetrators then stole a company vehicle to remove the stolen gold buttons and sludge from the site. Unfortunately, this type of activity goes on regularly. But managements of most mining companies are reluctant to discuss theft scenarios. So information pertaining to the theft of precious metals seldom becomes a newsworthy item. An audit conducted at a mine site with a gravity circuit recommended that the gravity recovery area be shut down until adequate protection could be provided. Although it was not connected with the audit, it was necessary to shut down the gravity area for a pro¬longed period because of problems with the gravity table. In the two months that followed, gold production at the site increased by about 31 kg/month (1,000 oz/month). It is difficult to attribute all of this increase to the theft of concentrates. But there was a good chance that at least part of the increase was due to the fact that concentrates were being stolen from the gravity area.

Jan 1, 1998

By G. C. Van Bever, J. J. Zaluski

Introduction The Surface Mining Control and Reclamation Act (Public Law 9587) (hereinafter the "Act" or "SMCRA") passed by Congress in August 1977 represents a comprehensive federal scheme for controlling surface coal mining and the surface effects of underground mining through permitting requirements, performance guidelines and reclamation planning. While the provisions of the Act have been the subject of numerous legal challenges and court battles over the years, it is difficult to identify a more controversial program within the Act than the provisions for designating lands as unsuitable for surface coal mining operations. The lands unsuitable designation process provides for the acceptance and review of petitions submitted by citizens or organizations seeking to have specified land areas designated unsuitable for all or certain types of surface coal mining activities. In filing these petitions, the interested parties or petitioners are required to make allegations about potential adverse impacts on people or the environment and submit evidence supporting their allegations. In 30 U.S.C. § 1272, Congress provided that "[a]ny person having an interest which is or may be adversely affected shall have the right to petition ... to have an area designated as unsuitable for surface coal mining operations." Under the Act, an area can be designated as unsuitable where the mining operation will (1) be incompatible with existing state or local land use plans, (2) affect fragile or historic lands, (3) affect renewable resource lands where mining operations could result in substantial loss or reduction of long-range productivity, or (4) affect natural hazard lands where such operations could substantially endanger life and property. In enacting SMCRA, Congress mandated that each state establish a process to determine which, if any, lands within the state are unsuitable for all or certain types of surface mining operations. In response to this federal legislation, the Kentucky General Assembly adopted a state regulatory program for surface mining that included provisions direct¬ing the Secretary of the Natural Resources and Environmental Protection Cabinet to establish a program for designating lands as unsuitable for surface mining as required by the Act. In recent litigation in Kentucky, several environmental groups filed a lands unsuitable petition, later joined by the University of Kentucky, challenging a proposal by Arch Mineral Corporation to surface mine over 3 million tons of recoverable coal. The petition sought to designate over 10,000 acres of land adjacent to Arch's proposed operations as unsuitable for surface mining operations, basically alleging that the mining would disturb an outdoor laboratory. The filing of the petition activated Kentucky's regulatory scheme for reviewing lands unsuitable petitions that can result in an absolute prohibition against surface mining on the petitioned land for historical, environmental and other related reasons. The designation process involves vague petition requirements creating a situation that Arch argued is devoid of constitutional due process and subject to abuse by the petitioner on many fronts. Arch maintained that the lands unsuitable regulations do not grant adequate protection to Arch's legitimate property rights under the due process clauses of the United States and Kentucky Constitutions and are thus void and unenforceable. The entire process resulting in a decision on the petition took just under 12 months in the Arch case, and although Arch was ultimately successful in preserving its right to mine, Arch's surface mining permit was held up for this period of time. This delay led to the cessation of mining operations by Arch and the idling of over 250 workers. This paper will review the lands unsuitable designation process and the significant implications the process has for existing surface mining operations, currently proposed operations and even those long-range operations not yet contemplated. Special emphasis will be given to Kentucky's lands unsuitable program. Finally, the recent litigation involving Arch Mineral Corporation and its effort to surface mine 81.5 acres of Arch controlled property will be utilized to illustrate this very unusual regulatory scheme. Regulatory background Chapter 30, Subchapter F of the Code of Federal Regulations (C.F.R.) promulgated to implement the provisions of SMCRA, requires that each state establish procedures under the state's surface mining program for designating non-federal and non-Indian state lands as unsuitable for all or certain types of surface coal mining operations. 30 C.F.R. § 764.1. The C.F.R. establishes minimum standards for state lands unsuitable programs and sets out requirements for filing a Lands Unsuitable Petition (hereinafter "LUP"), processing LUPs, decision-making guidelines and hearing requirements. Kentucky has adopted regulations providing for the implementation of the lands unsuitable process as part of the state's regulatory program under SMCRA. The following discussion summarizes the principle components of the Kentucky lands unsuitable program.

Jan 1, 1993

By Robert B. Fulton

The objective of this chapter is to discuss the interrelationship between industrial minerals and chemical manufacturing. It is intended to supplement rather than duplicate the commodity chapters. Particular emphasis is given to the pertinent chemical element and to market factors. Condensing this broad subject into a few pages of this handbook permits treating only the most important elements derived from industrial minerals. Hydrocarbons, which quantitatively dominate as raw materials for the chemical industry, are omitted, as are the metallic elements and the minerals covered in other "use" chapters such as phosphorous, potassium, and nitrogen for fertilizers, and titanium dioxide for pigments. The remaining six elements of major importance are: boron, bromine, chlorine, fluorine, sodium, and sulfur. These elements are treated individually under separate headings. [Table 1] affords an overview of the main industrial minerals, the chemical products derived from them, and end uses of the products. Salt brines have particular importance as raw material sources for the chemical industry. Table 2 is a chart of the chemical compounds derived from four types of brines: (1) Owens Lake-type brines, which are sources of boron and sodium compounds; (2) Midland-type brines, from which bromine, iodine, and chlorides of calcium, magnesium, potassium, and sodium are derived; (3) Searles Lake-type brines, yielding boron, bromine, lithium, magnesium, potassium, and sodium compounds; and (4) Silver Peak- type brines, produced mainly for lithium. MARKET ATTRIBUTES Some of the important market traits common to industrial minerals used by the chemical industry are: 1. They are international commodities, such as fluorspar and sulfur, which largely move to foreign consumers. 2. Grade, and freedom from deleterious elements are important factors affecting their usability in chemical processes. An example is salt (NaCl) used in electrolysis where ultrapure evaporated salt is required to meet rigid specifications. 3. Purified products take on the characteristics of specialty items and command a distinctly higher price than the basic commodity from which they are derived. 4. In practically all cases, chemical users require some sort of cleaning or beneficiation of the naturally-occurring mineral to bring it to specification, and individual specifications may vary from user to user for essentially the same use. 5. In some instances it is necessary to strike a balance between what the vendor can supply and what the buyer requires, with the result that specifications have to be eased to afford the needed materials in marginal cases. 6. Because they tend to be bulk commodities, low cost for handling and transportation are important and such costs may limit the area from which a chemical user can draw his supply. 7. Shipments are usually in bulk and frequently in multiple-car, full-trainload or full-shipload lots to reduce transport costs, which in turn may require large terminal investment facilities. 8. Purchases are generally by contract of one year or longer term, with spot buying playing only a minor role. 9. Contract prices are usually fixed in short term commitments, but may vary according to assay, with premiums and penalties for content above or below the norm; however, general practice is for specifications to be fixed in the contract with minimums being set for the desired material and maximums for undesired elements. In longer term contracts, prices are often escalated on labor, fuel, and other vendor processing costs. 10. Suppliers of individual commodities to the chemical industry tend to be limited in number and are generally medium- to large-size producers that supply a few major consumers. 11. The bulk of the mineral volume is for basic chemical uses, sulfur suppliers to sulfuric acid producers and fluorspar for hydrofluoric acid producers being typical examples. These basic chemical products then are used for the production of other products. 12. Shortage of a supply of adequate quality leads consumers to seek substitutes. In the case of fluorspar, much work is being done on recovery of fluorine from phosphate rock. Success in the form of fluorosilicic acid and/or hydrofluoric acid production could, in time, affect the hydrofluoric acid chemical industry. 13. Markets tend to be characterized by cycles of shortage followed by oversupply, with attendant wide price fluctuations. 14. Baniers to trade can have an adverse effect on the necessary movement of industrial minerals used by the chemical industry in international trade. Antidumping laws, quotas, and tariffs can disrupt or dislocate normal markets. 15. Chemical industry consumers may back-integrate for security of supply or for favorable economics, sometimes by joint ownership and often with experienced mining partners.

Jan 1, 1994

By D. J. Brunner, S. Mathur, D. McKinney

With the advent of faster micro-processors, the use of numerical methods to simulate complex fluid dynamic phenomena in three dimensions for use in design has become prevalent in the automotive, and turbo-machinery industries. The Computational Fluid Dynamics (CFD) method divides the region of interest into small control volumes which form the mesh representing the physical characteristics of the problem, and uses the finite volume method to intergrate the equations for the conservation of mass, momentum, energy and species over each control volume. Recent developments in CFD software expedite mesh generation, and enable the use of unstructured grids, comprised of tetrahedral volumes in three dimensions and triangular areas in two. CFD more accurately represents complex geometries and allows for relative movement of meshes enabling simulation of multiple moving bodies. 'ibis paper presents two examples of how CFD simulation can be used to assess mine and tunnel ventilation problems formerly addressed by application of analytical solutions which were developed assuming ideal incompressible conditions. CFD simulation is used to evaluate the impact of varying the airflow in a descentionally ventilated airway on the layering along the roof of smoke and hot gases resulting from a vehicle fire. Control of the smoke layer is required to enable safe egress from the vehicle, particularly if the vehicle is for personnel transport, and to ensure control of the fire contaminants throughout the ventilation system. The airflow required to prevent layering against the ventilation direction, calculated from the Bakke and Leach relations (Bakke and Leach, 1962), is compared with the CFD simulation results. An evaluation of the pressures, generated as a vehicle enters a tunnel portal, using CFD simulation, is also presented for unflared and flared portal configurations. These simulation results are compared with predictions derived using an analytical method which assumes one-dimensional and incompressible flow. Results of the CFD simulation are presented in an animated video format. SIMULATION OF BACKLAYERING In designing a ventilation system for a transit tunnel, the ability of the ventilating air to control and prevent backlayering of smoke and hot gases resulting from a vehicle fire is of prime concern. The buoyant nature of hot smoke causes it to rise relative to the colder, fresh air provided by the ventilation system. If the vehicle fire occurs in a descentionally ventilated tunnel, the smoke may tend to move upgrade in a layer above the incoming ventilation airflow. The layer may become thick enough to engulf a substatntial part of the tunnel cross-section upgrade of the incident that comprises the evacuation route. This effect is termed "backlayering' and it is similar to the development of methane layers in mines for which most studies related to backlayering have been done. Prediction Techniques Analytical A number of studies have been conducted (Bakke and Leach, 1962) to define the characteristics of this phenomena and as a result have produced relations which are used both in the mine and transit ventilation fields to define the air velocities required to control layering. In the transit industry the air velocity required to prevent the backlayering phenomena from occuring during a vehicle fire is called the "critical velocity" (Associated Engineers, 1975) and is dependent upon a number of factors: tunnel height, cross-sectional area and grade; ambient air temperature and density; and the heat release rate of the fire. Common practice in transit ventilation design is to provide an airflow which meets or exceeds the critical velocity. In order to determine whether or not the critical velocity can be achieved with a particular ventilation system, a one-dimensional simulation of the tunnel network is typically performed using programs such as the Subway Environment Simulation program (SES) originally developed in the late 1970's (Associated Engineers, 1980). The results obtained from SES are compared to the critical velocity to determine the adequacy of the ventilation system. Computational Fluid Dynamics For the backlayering simulations, a commercial CFD code which has been used successfully in a wide variety of engineering applications, was used. It provides numerous options for modeling laminar and turbulent flows, multiple turbulence models, definition of multiple species and chemical reactions between them, a variety of boundary conditions (including constant pressure and constant velocity inlets) and the ability to apply user-defined FORTRAN subroutines. It includes the ability to model conductive, convective, and radiative heat transfer. FLUENT also permits the use of "body-fitted coordinates" to match the computational mesh or grid to complex real-world geometries. Computational Fluid Dynamics Model The model developed to simulate the backlayering phenomena is comprised of an airway of rectangular cross-section, 4 meters wide, 4.5 meters high, and 200 meters long. A laterally

Jan 1, 1995

By R. W. Carn, D. Vincent

During the next 15 years, US coal production is expected to double, with the increased production evenly divided between the East and the West. Along with greater production, coal export markets should increase dramatically from East, West, and Gulf Coast ports. The annual overseas export capacity of US coal-loading terminals is expected to rise from 147.1 Mt (162.1 million st) in 1981 to a minimum of 278.1 Mt (306.6 million st) in 1985, according to the US Maritime Administration. Increased coal production and use will lead to more development of import and export terminals, a vital link in the coal transportation chain. With continually escalating capital costs and the competitive markets that the terminals will serve, a well designed and efficient terminal is necessary. This article begins a two-part series that presents concepts presently used in coal export terminal design. Part I looks at site selection factors and equipment needs, while Part II will examine environmental considerations in building a terminal as well as typical capital and operating costs. The world is nearing the end of the oil era. In a few years oil will not be available to sustain the growth rate and increasing standard of living we have known in our lifetime. The big question is what energy era are we moving into? With the decline of readily available oil reserves and rapidly increasing prices, many countries are trying to switch to alternate energy forms. While intensive efforts to find new oil reserves continue, alternate energy sources such as natural gas, coal, synthetic fuels, nuclear, hydroelectric, solar, and wind power are being developed. Recent indications are that coal is expected to bridge the energy gap over the next 25-30 years until the technology and economics of the alternate energy forms reach satisfactory levels. Use of coal for energy is receiving strong attention due to its long-term availability (200-300 years minimum), relative ease of development, and its low cost per unit of power produced. By the year 2000, it is expected that 25% of world energy supply will be met by direct coal combustion and possibly another 5-10% by synthetic fuel from coal. Coal's expanding share in the world energy market, along with an increase in coking coal requirements, will result in a large increase in the world's seaborne coal trade. Recent statistics and projections for the future are shown in Table 1. This phenomenal development rate includes increases in both coking and thermal coal requirements. Because of the rapid increase of seaborne coal trade during the last 10 years and the even greater projected increase of trade to 2000, various sectors of the coal industry are faced with enormous technical challenges and huge investments in equipment, land, transportation systems, and port facilities. Very large bulk terminals are under development throughout the world. Latest surveys indicate that there are about 30 new coal export and import terminals under consideration and at least 30 existing terminals have expansion programs planned or underway. With the high cost of borrowed capital and rapid inflation rates there is great emphasis on new planning and design techniques to minimize capital and operating costs of coal transportation systems. Terminals A total coal supply system can be considered to consist of one or more mines; a train, barge, truck, or other haulage system; an export terminal; a fleet of bulk carriers; a receiving terminal; and possibly, local inland distribution networks that include barges and railways. Terminals, though only a small link in the total transportation system, play a key role in overall system efficiency. At ports or inland distribution centers, terminals act as transportation links bringing trains, ships, barges, or trucks together for cargo transfer and temporary storage. A well-designed terminal can provide maximum independence between two modes of transportation and optimum freedom for intermodal interference. A terminal acts as a buffer between the two transportation modes by providing sufficient storage capacity so a ship need not wait for its cargo on, for example, a train-by-train basis, but can load immediately from the ready stock. Similarly, a train need not wait for a ship to unload its contents but can dump immediately into storage. A terminal also can be used to properly mix various types of coal to satisfy a buyer's requirements. Consider the relative value of various production and transport segments for a typical steam coal

Jan 6, 1983

By E. D. Attanasi, J. H. DeYoung

Several factors contributed to continued declines in mineral-exploration activity in the US in 1985. Low metal prices and, what appears to be worldwide chronic excess capacity in copper, molybdenum, lead, and uranium, have resulted in mineral-exploration expenditures remaining anemic. Economic recovery could result in a healthier mining industry and more cash flow to fund exploration. This is because general economic activity and US mining industry activity have historically been closely linked. However, as the worldwide economic recovery has expanded, the mining sector has continued its downward slide. New cuts in industry exploration budgets in 1985 shocked those who thought the exploration situation could not become worse. Some personnel and equipment had been redirected from base metals exploration to precious metals in the past few years. Last year, continued reductions in exploration sent many professionals out of the mining industry. Recent staff reductions or consolidations of operations were made by Noranda, Chevron, Molycorp, and other exploration companies. The latest data from the Society of Economic Geologists (SEG) summary of exploration statistics show that professional staff at year end in major US exploration companies (domestic and foreign operations) fell from 2355 in 1981 to 1868 in 1983 and 1277 in 1984. By the end of 1985, two economic trends were established that could improve the future profitability of mining and hence exploration. First, the price of crude oil began a decline. If sharply reduced energy prices increase worldwide economic expansion, the substantial excess capacity in some of the base metals industries could disappear, and prices could improve. Furthermore, if energy price declines reduce mining and processing costs significantly, metals may recapture some lost markets. The decline in oil revenues has already encouraged some oil-producing countries, such as Venezuela, to look toward development of mineral resources to earn foreign exchange for debt repayment. Second, the decline of the dollar by 21% during 1985 could also help US producers meet foreign competition. During 1985, industry restructuring continued as many oil companies sold off mining subsidiaries and minerals properties. Gold, silver in new discoveries Precious metals continued to dominate the announcement of new discoveries and exploration projects in 1985. A review of domestic exploration and development activities reported in several industry journals shows that 60% to 80% of these projects were directed primarily at precious metals, particularly gold. Base metals exploration activities frequently involved polymetallic deposits with gold or silver values. Because much of this exploration was done on identified targets (on-property exploration), the decrease in wildcat or grassroots (off-property) exploration may be more substantial than indicated by reductions in total exploration activity. Significant gold discoveries in 1985 included several in Nevada, among them the Genesis property of Newmont (near the Carlin mine), Goldfields' discovery of the Chimney deposit in Humboldt Co., and Freeport's discovery of two mineralized sites near Jerritt Canyon. Gold exploration continued to be focused in the western US and Alaska, but gold production starts at the Haile mine in South Carolina, and the Ropes mine in Michigan as well as Amselco's feasibility studies on deposits near Ridgeway, SC, are evidence that gold exploration is not limited to the West. The dominance of gold projects in exploration is not limited to the US, as demonstrated by gold dis¬coveries and exploration projects in Australia, Brazil, Canada, the Caribbean region, China, Guinea, Ivory Coast, South Africa, the South Pacific islands, and Thailand. From the standpoint of US metal miners, it is perplexing that worldwide exploration and development is also taking place in copper, zinc, tungsten, and other metals with depressed prices. During 1985, the US Geological Survey's efforts to map the sea floor of the Exclusive Economic Zone shifted from the Pacific Coast to the deep water areas of the Gulf of Mexico and to areas off the coast of Puerto Rico and the Virgin Islands. An atlas containing sea-floor maps of the west coast area was published as US Geological Survey Miscellaneous Investigations Series Map 1-1792. Results of the 1985 surveys are expected to be published by January 1987. Exploration trends - Statistical evidence Data from the SEG showed continued decline in the US mining industry's exploration expenditures through 1984. The share of US companies' domestic exploration expenditures directed toward base and precious metals has increased from 51% to 84% from 1980 to 1983 and to 86% in 1984. US mining companies spent about $0.67 of each exploration dollar in 1984 in the US. However, this represents an increase from earlier years. The 1983 data also show that firms spending more than $5 million on exploration accounted for 77% of exploration expenditures. Since 1981, the Bureau of Land Management (BLM) has been assembling data on claims and an-

Jan 5, 1986

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

Hydrodynamic investigations in exploratory boreholes and grouting holes are conducted for the purpose of obtain¬ing information about the hydraulic properties of the hydrostratigraphic section to be intersected by the proposed underground workings. The information obtained from the investigations provides the basis for calculating the hydrau¬lic coefficients of fractured permeable rock, the dimensions of the anticipated grout isolation curtain(s) around the un¬derground workings, the number and location of grouting holes, the injection pressure modes, and also the volume(s) of grout that will be required (Anon., 1976, 1978). The following data on each aquifer are obtained from the investigations conducted in monitoring and grouting bore¬holes and the analysis of the results: 1) the top of each hydrostratigraphic unit, 2) the thickness of each unit, 3) the ground water fluid potential distribution in each unit, 4) the coefficient of permeability, 5) the piezoconductivity, 6) the fracture porosity, 7) the geometry of the fractures in the rock, 8) the elasticity-compressibility coefficient of the fractured rock, 9) the chemical composition of the ground water, 10) the direction of flow of the ground water, and 11) the expected inflow rate of water into the shaft, drift or tunnel. STG uses its DAU-3M type flowmeter to conduct in¬vestigations of directions of flow in vertical, inclined and horizontal drillholes. The DAU-6 instrument is used to de¬termine the direction of flow of ground water in each frac¬ture or fractured aquifer. Various singular and double DAU type packers are used for pumping and for injection studies (tests) and for flowmeter investigations. Normally the instruments enumerated above permit in¬vestigations to be conducted in each separate aquifer with¬out reinforcing the holes with casings. On the basis of these investigative data, both the hydraulic properties of unfractured rock and the hydraulic properties of the fractured rock are estimated. Dual porosity rocks require special attention because they tend to segregate the grout. 3.1 FLOWMETER INVESTIGATIONS IN BOREHOLES The STG flowmetric methodology is based on the mea¬surement of the ground water flow rate through the borehole by hydrostratigraphic interval after the disturbance of the hydrostatic equilibrium in the "hole-aquifer system" (after pumping or injecting). The relationship of the head changes to the discharge into or from a particular hydrostratigraphic unit obtained during the tests serve as the basis for calcu¬lating the hydraulic properties. Flowmetric investigations facilitate the determination of the number of aquifers, their depths, their thickness, the hydraulic properties of the fractured rock and the magnitude and direction of the flow of ground water. 3.1.1 FLOWMETER HARDWARE STG conducts flowmetric investigations in boreholes using its DAU-3M-108, DAU-3M-73, DAU-3M-57 and DAU-3M-44 instruments.' They have respective external diameters of 108, 73, 57 and 44 mm. The type of flowmeter selected for use depends on the borehole geometry and the technological scheme for carrying out the investigations. Boreholes with a drilling diameter of 76-93 mm are inves¬tigated with the DAU-3M-73 flowmeter; boreholes drilled by bits with a diameter of 112 mm and more are investi¬gated using the DAU-3M-108 flowmeter. The DAU-3M¬108 and DAU-3M-57 instruments are used for flowmetric investigations with a packer. 3.1.1.1 The Downhole Sensor The sensor design of the DAU-3M-73 hole flowmeter is shown in Fig. 2. The design of the DAU-3M-108 instru¬ment is similar to the design of the DAU-3M-73 instrument. The frame of the flowmeter sensor shown in Fig. 2 consists of a casing, an upper and lower centering mount and two rings to which the guiding rods are attached. The upper rods are built into the connector bushing; the lower rods are built into the coupling sleeve. The borehole cable is attached using a half-coupling, a packing ring and a constriction nut. Thus, the frame of the flowmeter sensor is made so that the free passage of water to the impeller is facilitated along with the necessary rigidity. The primary moving component of the flowmeter is the double-bladed impeller, which rotates on cobalt-tungsten pivots and agate thrust bearings. Special extended air cham¬bers protect the supports of the impeller from the action of the borehole fluid which may contain fibrous and abrasive particles. The air located in the chambers shields the sup¬ports from direct contact with the borehole fluid when the sensor operates in a borehole. The hollow casing of the impeller serves the function of a lower cap. The upper cap is attached to the casing using a threaded connector; it is affixed also with a lock-nut. An adjusting screw with a

Jan 1, 1993

By M. J. Howes, C. A. Nixon

INTRODUCTION A safe heat stress control strategy for an underground mine has three elements: Application of an environmental measure which reflects physiological strain with sufficient accuracy for the range of conditions encountered underground. Acceptance of a functional relationship between the environ- mental measure and human performance which is used to optimise the environmental conditions achievable with either ventilation or ventilation and refrigeration. A management control strategy based on the environmental measure which is designed to ensure that work in environments where excessive physiological strain may occur is prevented and corrective action is initiated. The environmental measure that reflects physiological strain is the link between the three elements and, since the turn of the century, the discussion of the merits of various indices has been prolific. One problem in selecting a suitable measure or index is the ease with which it can be physically obtained relative to accurately reflecting the physiological strain. For example, wet bulb temperature is simple to measure and, for a particular mining sys- tem, it may adequately represent physiological strain, however, it would not necessarily provide the same relatively safe measure in a different mining system. The acceptance of a measure which can be universally applied has been confounded by both development and predisposition. That is not to say that there is only one "correct" measure and all others are unsuitable. It is self evident that if the application of a particular index has resulted in adequate control, then that mea- sure is correct for that situation. However, an understanding of the limitations is necessary to ensure that adequate control is maintained as mining conditions change. Almost 100 years after the question of heat stress in mines started to be dealt with in a collective manner, an analysis of the available information is leading towards a general strategy to control this problem. In the paper, the developments in heat stress assessment are briefly examined and followed since the earliest published observations on the effect of heat in mines (Haldane, 1905), efforts to determine a relationship between an environmental measure and human performance are reviewed and summarised and the benefits of control strategies such as acclimatisation and shortened shifts are discussed as they relate to Mount Isa Mines. The results of testing the prototype air cooling power instrument are discussed and a heat stress control strategy outlined. HEAT STRESS AND AIR COOLING POWER The operation of the human engine is analogous to other engines where the conversion of chemical energy from the oxidation of fuel to useful mechanical energy is not 100% efficient. In a diesel engine it is about 33% and in a human engine less than 20% resulting in at least five times as much heat produced by the meta- bolic process as useful work done. Metabolic energy production is related to the rate at which oxygen is consumed and is about 340 W for each litre of oxygen per minute. Using measured oxygen consumption and an average body surface area of 2.0 m2, the approximate metabolic energy production associated with different mining tasks is (Morrison et al. 1968):- • Rest, 50 W/m2 • Light work, 75 to 125 W/m2 (machine, LHD or drill jumbo operators) • Medium work, 125 to 175 W/m2 (airleg drilling, light construction work) • Hard work, 175 to 275 W/m2 (barring down, building bulkheads and timbering) • Very hard work, over 275 W/m2 (shovelling rock) Heat balance is achieved when the rate of producing heat (the metabolic heat production rate) is equal to the rate at which the body can reject heat mainly through radiation, convection and evaporation. Heat exchange between the lungs and the air in- haled and exhaled is normally less than 5% of the total and there- fore usually ignored. Any heat not rejected to the surroundings will cause an increase in body core temperature. Since heat stress is related to the balance between the body and the surrounding thermal environment, the main parameters required to be known when determining acceptable conditions are those associated with the heat production and transfer mechanisms. These can be summarised as follows: Metabolic heat production rates (M - W) Skin surface area (A3) (and effects of clothing) Dry bulb temperature (t[ ]) Radiant temperature (t[ ]) Air velocity (V) Air pressure (P) Air vapour pressure (e [ ]) The rate of heat transfer to or from the environment depends on the equilibrium skin temperature t, and the sweat rate S,. These in turn depend on the response of the body to the imposed heat stress and the effect of thermoregulation (Stewart, 1981). Thermoregulation The body contains temperature sensitive structures which send impulses to the brain at a rate depending on the temperature. Both hot and cold signals can be differentiated and the thermoregulatory response ahivated according to which signal pre- dominates. If "cold" signals are dominant, body heat loss is re-

Jan 1, 1997

By Austin Whillier

INTRODUCTION Use of air-cooling or refrigeration equipment in underground mines is needed when conventional ventila¬tion techniques do not maintain acceptable environ¬mental temperatures in working areas. Because refrig¬eration can be very expensive, it should be implemented only after all possible and practical steps have been taken to eliminate or reduce heat sources in the mine. As an example, to prevent the main ventilation fans from contributing heat to mine air, they should be located in the air return and not in the air intake. It is particularly important to prevent any direct contact between hot water and ventilation air, especially in mines which encounter large flows of hot fissure water. Any water hotter than the prevailing wet-bulb temperature of the ambient air must be removed by pipes located as close as possible to the water source. This hot water must not be allowed free contact with the incoming ventilation air at any time during the water's passage out of the mine. Although insulation of the pipes carrying the hot water is seldom necessary, direct contact between the air and the water must be prevented so the warm water cannot evaporate. REVIEW OF COOLING PRACTICES Spot Coolers vs. Centralized Refrigeration To eliminate a few specific hot places in an otherwise cool mine, it is possible to use devices known as "spot coolers." A typical spot cooler that uses chilled water is shown in Fig. 1. These devices consist of self-contained refrigeration units that are often mounted on rail cars for haulage to hot spots. The cooling capacities of such spot coolers usually are limited to about 100 kW or 30 "refrigeration tons." A refrigeration ton represents a cooling rate that produces 1.0 st of ice in 24 hr; that is a cooling rate of 3.517 kW (200 Btu per min). Typically, the electric-power consumption to drive the compressor motor of the refrigeration plant in mines is 1.0 kW per refrigeration ton, corresponding to a coefficient of per¬formance of about 3.5. The principal difference between spot coolers and centralized refrigeration plants is the method of re¬jecting heat from the refrigeration system. Centralized refrigeration plants always discharge heat into the reject or return airflow of the mine; often that is the primary influence in selecting the location for the underground refrigeration plant. Heat from spot coolers usually is rejected into drain water or into air that is not flowing to the location requiring the cooling. As a result, spot coolers remove heat from troublesome hot spots in the mine, injecting that heat-plus the electrical energy used by the cooling unit itself-into other working areas where the ambient conditions are cooler. In effect, this is "robbing Peter to pay Paul." In deep, extensive mines, spot coolers usually pro¬vide only temporary and, over the long term, expensive solutions to localized cooling problems. Centralized re¬frigeration plants are preferred for such mines, with cooling distributed throughout the mine as required. Fig. 2 illustrates a typical underground centralized re¬frigeration plant. Centralized plants lend themselves to improved maintenance at reduced costs while offering the economy of size. Refrigeration plants of larger unit sizes have considerably lower initial costs than smaller unit sizes. The remainder of this chapter is devoted to large refrigeration plants, with no further consideration of spot coolers. Cost of Refrigeration Total Cost: The total cost of refrigeration amounts to about $200/kW of cooling per year (1981 US $). This total cost breaks down into approximately three equal parts: 1) Financial charges, which include the interest and amortization on the capital cost of the initial installation, and the cost of necessary underground excavations. 2) Operating and maintenance costs which include the cost of the electric power to drive the refrigeration plant's compressors. 3) Distribution costs which include costs for pump¬ing, insulated piping, and air-to-water heat exchangers. The local cost of electric power, the number of operating months per year, and the method of refrigera¬tion distribution all contribute to the actual costs in¬curred in a given application. However, the variations usually are limited to no more than ±30% of the $200/ kW per year total cost figure. Cost Per Ton: Refrigeration cost per ton of mineral production can be calculated if the annual production tonnage from the refrigerated section of the mine is known. In most cases, this cost will be less than $1 .00/t. However, in deep mines with high rock temperatures, such as those found in South Africa, the total cost of refrigeration can increase to several dollars per ton of broken rock. Distribution In deep extensive mines, distributing refrigeration often accounts for about half the total cooling costs. As a result, careful consideration and planning must be

Jan 1, 1982

By A. J. Richardson

Before the development of the underground stoping and mining costs can be considered, certain facts about the ore body, the proposed mine, markets, etc., must be known or determined. In the case to be studied, the zinc-lead mineralization occurred with a narrow vertically dipping structure of undetermined length and vertical extent. Exploration completed to date has revealed 6.5 mil¬lion st t of proven reserves. A further 820,000 st of in¬dicated reserves has been outlined and this tonnage is considered capable of being expanded by a factor of approximately four after more detailed drilling. After studying the market conditions and completing a very preliminary feasibility study, it was decided that production would be 730,000 stpy (or 2000 stpd) of ore. First year production would be at the rate of 1500 stpd. The main design criteria for the selection of the min¬ing methods are minimizing surface subsidence, maxi¬mum recovery of the ore body, maximum degree of grade control, maximum productivity, and safe working conditions. Two basic extraction systems are considered capable of meeting these requirements: mechanized cut¬and-fill stoping and sublevel long-hole stoping with filling. The primary development system of the mine has been designed to give maximum flexibility in stoping systems and layout and to permit changes if considered necessary as a consequence of actual production ex¬perience. At the present time, access to the mine is by a circu¬lar concrete lined vertical shaft, 16 ft diam, sunk to a depth of 1380 ft. Two exploration levels have been driven within the ore zone at depths of 165 and 1246 ft below the surface outcrop. The development to date had the objective of sampling the mineralization and produc¬ing detailed information on the outline of the ore body and the distribution and controls of zinc and lead values. In an attempt to satisfy the basic design criteria for the mine, it was decided that production would be best achieved by a combination of 40% sublevel long-hole stoping and 60% cut-and-fill mining. Costs of exploration and capital development of per¬manent underground facilities are normally written off over the life of a mine. Production expenditures, on the other hand, are of a temporary nature and are normally charged as and when incurred as an operating expense. Reasonably accurate predictions of mine production costs can be built up from engineering design and estimates of individual mine activities for ultimate inclusion in the comprehen¬sive data required for financial decision making. The simulated operations can be costed on a detailed basis in the form of a monthly operating budget. The budget format can be generalized or detailed, depending upon the scope of the project. However, ex¬perience suggests that a fairly detailed format has the advantage of assuring that all significant cost items are included. For underground costing it is suggested that the budget structure include five major cost centers (i.e. development, diamond drilling, ore extraction, hoisting/ transportation, and general mine expense). These in turn are detailed under numerous subheadings. The mechanism for compiling an operating budget will be illustrated. Because of its relative simplicity, ore extraction under sublevel long-hole stoping has been chosen for illustration. All other activities, simple or complex, can be estimated in similar fashion. BLOCK AND STOPE DEVELOPMENT Long-hole blocks, used where advantageous, will be up to 250 ft in height, depending upon the vertical con¬tinuity of the mineralization, and approximately 300 ft long. Drawpoints will be at 36-ft intervals and serviced by loading crosscuts driven from a footwall drift parallel to and close to the ore zone. Pillars between the stopes will be 50 ft wide. Stopes will be drilled off with vertical rings of blastholes drilled from sublevels approximately 60 ft apart vertically. This drilling will be done by percussion drilling machines (31/2 in.) mounted on a trackless drilling rig. Load¬haul-dump (LHD) equipment will be used to move broken ore from the drawpoints to the orepass connecting to rail haulage systems. On completion, long-hole stopes will be backfilled to prevent caving and to facili¬tate later pillar removal. From a planned stope layout, a forecast of produc¬tion and development is made in Table 1. Table 1. Block Tonnage and Stope Development Quantity Ore Waste Total ore block 375,000 st 2 stopes 310,000 st 1 pillar 65,000 st Access crosscuts, 4 at 100 ft 400 ft Drill sublevel drifts, 6 at 300 ft 1800 ft Stope raises, 3 at 250 ft 750 ft Undercut sublevel drifts, 2 at 300 ft 600 ft Loadout crosscuts at 35-ft intervals 550 ft 100 ft 3300 ft 500 ft Total development footage 3800 ft Tons per ft of development 987 st

Jan 1, 1982